l1nited States Patent Office
1
3,777,004
PROCESS FOR HEAP LEACHING ORES
Arthur W. Lankenau and James L. Lake, Lakewood,
Colo., assignors to Hazen Research, Inc., Golden, Colo.
No Drawing. Filed May '10, 1971, Ser. No. 141,960
Int. C1. BOld 11/00; COlg43/00
U.S.' CI. 423-20 26 Claims
ABSTRACT OF THE DISCLOSURE
A method for recovermg metal values from hirge deposits
of low grade ores or from high grade ores existing
in small ore bodies in an outdoor operation near the mine
which comprises making a pile of the ore and successively
leaching selected depths of the ore from the top of the
pile down by adding leaching agent and sufficient moisture
to leach the metal contained in each selected depth with
intermittent stirring of the selected depth of ore until
the metal is solubilized followed by removal of the leached
section from the pile, separation of the leach solution and
solid tailings and recovery of the dissolved metal values
from the leach solution.
BACKGROUND OF TIlE INVENTION
There are large existing bodies of ore of such low
grade that it is not economical to mine and refine it to
recover the metal values in it by conventional recovery
techniques. Examples are low grade uranium, copper and
vanadium ores, and others. There are also small are bodies
of high grade ores of these and other metals which caunot
be economically processed because the total amount of
metal available is too small.
The cost of refining these ores includes capital investment
and maintenance of processing equipment and buildings,
as well as the labor' involved. A further cost is
the transportation of are in bulk from the mine to the
processing site. These and related expenses make it economically
prohibitive to refine the ores by conventional
means. As the higher grade ores in large are bodies reach
the point of exhaustion, attention is being focused on the
development of economical methods for the processing and
refinement of large low grade are bodies.
Another disadvantage attendant to present metal recoverytechniques
used for certain ores is the resultant
pollution problems in congested areas resulting from byproduct
fumes' and gases.
The present invention has as its principal objective a
method for processing and refining both low grade and
high· grade' ores in the open near the mine site away
from congested areas and with a reduction of capital
investment, housing, labor, equipment, transportation and
other costs.
In accordance with the invention the are in the mine
or in piles on the ground made after crushing is leached
in situ in the open air by applying moisture as required
and leaching agent to a selected depth of ore accompanied
by intermitting stirring of the selected depth of
ore until leaching is complete for the selected layer of
ore. If the ore contains sufficient moisture none is added.
The layer of leached are in a slurry is then transferred
to a. tank or basin for separation of the pregnant leach
splutionfrom the tailings and metal values recovered from
the leach solution by ion excange solvent extraction or
other conventional means. By "leaching in situ" is meant
leaching in the mine without first removing the are or
leaching the are outside the mine in a pile. The order of
addition of water and leaching agent is not critical. Water
may'pe added bef()re or after the leaching agent or water
and leaching agent may be added together.
Successive layers of ore are leached in the above man-
3,,777,004
Patented Dec. 4, 1973
2
ner until all of the are has been leached. The raffinate
from the ion exchange step is reused in the leaching
operation. Multiple recovery operations can be conducted
simultaneously with continuous use of settling basin and
5 ion exchange equipment, with are from the leach solution
from one pile being recovered while are from another
pile is being leached.
It has been found preferable to use a depth of top
section of are about 6-12 inches for each successive
10 leaching operation. A laboratory analysis is first made
of the are by the use of grab samples to determine the
approximate ratio of leaching agent to are necessary.
Adequate water must be present to insure optimum
leaching conditions. Water must be added if the are lacks
15 sufficient moisture or if the leaching solution is not sufficiently
dilute to supply the necessary water.
. Stirring the section of the pile being leached is highly
important to the process. This insures that the leaching
solution does not percolate below the top section being
20 leached and, more importantly, it insures good contact
of the are with the leaching agent. Stirring can be accomplished
with conventional equipment, such as, a harrow,
cultivator type instrument, etc. This stirring step exposes
the are to air for oxidizations which is beneficial to the
25 recovery of metal values in a reduced state.
It is important that the proper amounts of moisture
and leaching agent are maintained in the are at all times
and that both be uniformly distributed throughout the are.
The approximate amount of leaching agent is ascertained
30 in advance by laboratory tests. The progress of leaching
can be determined by analysis of grab samples and additional
amounts of leaching agent and water added as necessary.
The necessity for additional moisture is apparent
from observation of the ore, the requirement being that
35 the are is maintained in a moist condition at all times.
Uniform distribution of moisture and leaching agent is
accomplished by stirring and the stirring must be done
intermittently as necessary while the are is being leached.
Otherwise, there will be non-uniform leaching through-
40 out the section and undesirable amounts of leaching agent
will percolate beneath the selected top section. Finally,
the stirring insures that the leaching operation is restricted
to the top section. After tests indicate the leaching is
complete the leached section to the required depth is
45 readily removed with a bulldozer, front-end loader, scraper
or similar device.
The preferred order of steps after the ore is in a pile
with a substantially flat top surface is to uniformly sprinkle
water over the top of the pile, stir the top to a depth of
50 about 6-12 inches, add leaching solution uniformly to the
top of the pile, stir the leaching solution uniformly into
the wetted top 6-12 inches of the pile, followed by intermittent
stirring of the ore and addition of water as required
to maintain the ore in a damp condition over a period
55 ranging from a few days to several weeks.
The leached are after removal from the pile is transferred
to a sluice box to form a slurry and the latter transferred
by pumping, gravity or otherwise, to a settling basin
which can be a hole in the ground or pond lined with
60 plastic to make an impermeable pond area.
After the solids in the slurry have settled, the clear
supernatant leach solution is transferred through piping
to an ion exchange station where metal values are re-
65 covered from it. Other recovery procedures may be used.
The raffinate from the ion exchange step is recycled to
the pile areas for use as a leaching agent or is sprayed
into the sluice box to pulp the leached ore.
The process was extensively applied to low grade ura70
nium ore from the Maybell, Colo., and Baggs, Wyo., area,
and to native copper and vanadium ores using concentrated
sulfuric acid leaching agent. The low grade are
3,777,004
EXAMPLE 1
The results set forth in the ". following table were obtained
on a uranium ore withcoricentrated sulfuric acid
leaching agent using the above-described procedure.
TAB:J:,E 1
As seen from the table, a .98 percent recovery was
obtained from the urarrlum, ore in load .1 containing only
0.10 percent uranium based on U30 8••Recoveries from
the ,five loads of loW grade ore, varieci ,from 84. to 9,8
percent.
Agitation leaching tests ,run, on. saDlples taken from
loads of ore leached in the pilot plantrunsshow~d·that
the percentage of ,leaching agent per. weight of. ore is
fairly accurately obtained in the agitatioI\ leaching tests.
For, example, agitation. leach ,tests for two representative
55 ASSay~ percent
,as U.Os, percent Acid added,
lb./ton
are Slulce Extracfeed
" feed, Tails.; tion .93.2% 100%
Load No.:
60
2L________________"_"___"__ 0.10 0.10 0.002 98.0 44 41.0
3____________" 0.16 0.12 0.003 98.1 37 34.5
4____________" 0.05 0.04 0.003 84. 0 48 44.7 0__________ ••_ 0.10 0.10 0.008 92.0 34 31. 7
0.29 0.20 ' 0.007 97.6 49 45.7
Average. __~;;;; ____.-;;;;;_'._.;_"._ 0.005 ~3.9 42.4 39.5
50
35
30
4'·,
picked up with a front-end loader. and dumped onto a
%" screen in a sluiCe box located adjacent a tailings
pond. The ore was washed. with a solution jet to form
an ore-solution slurry using a high pressure stream of
acid raffinate from the ion exchange recovery step which
was sprayed down into the sluice box to pulp the leached
ore. The pulp was flowed to the tailings pond through
a gravity launder. ·Thepond was lined with plastic sheeting.
A large excess of raffinate was used to keep the con-
10 centration of uranium in the leach liquor at a low level
(0.10-0.20 gram U30 8 per liter) to minimize the uranium
losses in the settled solids. The clear pregnant leach liquor
was transferred from the pond through a hose to the ion
exchange circuit. '
15 The ion exchange system was operated by passing the
uranium-bearing solution from the tailings pond upward
through the resin in the column. Each of the two feet
diameter columns was loaded with, ten cubic feet of
16 x 20 mesh Dowex 21K resin,' a strong base anion
20 exchange resin. Another suitable' resin is Rohm and
Haas Amberlite XE-270. Other conventionally used resins
are suitable. While one column was being loaded, the
other column was being eluted or· "stripped" to remove
the extracted uranium material from the resin. The
25 uranium was eluted from the loaded resin with one molar
sodium chloride and 0.15 normal sulfuric acid solution.
After elution, the resin was washed with water whiCh
was then transferred to the precipitati()n tank with the
pregnant stripping solution.
The uranium was precipitated' as yellow cake (U30 8)
by neutralizing with anhydrous ammonia at a pH of
about 7, in accordance with conventional procedures.
The precipitation efficiency was found to be 99.8 percent
from a relatively low grade elution solution.
The analysis of the yellow cake was near· commercial
specifications. It was found that use of a two-stage precipitation
method reduced the iron content to 0.12 percent.
In this method the pH is raised to 4.0 with ammonia,
the resulting iron precipitate. removed by filtra-
40 tion, and the neutralization and precipitation completed
with ammonia.
The sodium content of the .yellow cake was reduced
to 0.26 percent by agitating the concentrate· for 2 hours
with a 200 gram per liter ammonium sulphate solution
45 before filtering and washing. With two-stage precipitation
and ammonium sulphate washing,the yellow cake contained
only 0.05 percent chloride and 0.00 percent
calcium.
1 1
Bleed to Waste
1(Eluate) r NHa
1
1
1
Filter Press ----I
1
Filter Press --> Iron Product
Solution
Precipitation Tank
Precipitation Tank <- NHa
1
(Raffinate)
Sluice Box
1rH2S04
-1
,' (1) Crushed and blended by bnlldozer.
Treatment Pile (2) Concentrated acid sprinkled on
surface.
1
(3) Moistened and turned for 8 days:
(Scraper Haul) r Water
Sealed Tailings Pond ~ r-Storage Tank
IX Column
Low Grade are
Uranium Concentrate
Figure 1
Truck-load lots of selected ore samples were brought
to the pilot plant site from existing open pits and level
piles of the ore were made preparatory to leaching. Samples
from each lot were analyzed to determine the approximate
ratio of acid to ore necessary for leaching. The
soft sandstone ore was broken into small particles by
running a tractor over the pile in several directions. Concentrated
sulfuric acid was then sprinkled uniformly over
the top of the ore. A spring-toothed cultivator or harrow
was then used to stir the top section of the piles to a depth
of between 8-12 inches. The depth of the teeth of the
cultivator could be adjusted so that the depth of the section
being treated could also be regulated. Water was 65
then sprinkled uniformly over the tops of, the piles to keep
the upper layer moist and the ore was cultivated every
day or every other day nntil a total of eight days had
elapsed. The moisture content of the sections being treated
was maintained at about 10 percent. Water was added 70
periodically to maintain the moisture content, and the
heaps cultivated regularly. The same procedure was followed
using piles of ore on the ground about 8 inches
deep, 12 feet wide and 40 feet long.
At the completion of the leaching operation the ore was 75
3
deposits are located too far from operating mills to economically
transport them and recover metals therefrom
by conventional processes. As is the case with many other
low grade ores, the presently known reserves are believed
to be insufficient to economically support the construction 5
of a conventional ore processing plant at the site. The
principal objective is to provide a process requiring relatively
low capital investment and operating costs whiCh
produces a high purity metal product.
The process followed the following flow sheet:
FLOWSHEET FOR TREATING LOW GRADE
MAYBELL AND BAGGS URANIUM ORES
3,777,004
9.62
7.60
Lb.H2S0.
per lb. Cu
extracted
65.0
41. 2
0.28
0.47
Tails Percent Cu
percent Cu extractiou
TABLE 2
EXAMPLE 2
EXAMPLE 3
100
50
H2S0.,
lb./ton
The mineral used was a low grade copper oxide
mineral.
The top layer of a pile of ore for a depth of about 8"
was leached. Concentrated sulfuric acid was sprinkled
over the top of the ore and the ore stirred for a depth
of about 8". Enough water was sprinkled over the ore
to give approximately 10 percent moisture. The ore was
mixed daily and the moisture content maintained for approximately
ten days. At the end of ten days, the ore was
repulped with water for one-half hour, then filtered and
washed. The copper values were recovered from the leach
solution by standard recovery procedures.
The heads contained 0.80 percent copper, the ore being
ground to a 28 mesh size. The results obtained from two
55 tests are as follows:
60 Test No,: L •.•..._
2•••• • _
A vanadium bearing sandstone and clay low grade ore
was used. The heads contained 3.35 percent V205 and
the ore was ground to a -6 mesh size.
Again, the top layer of a pile of ore about 8" deep
was leached. Concentrated sulfuric acid was sprinkled
70 over the top of the ore and enough moisture added to
give about 1'0 percent water. The pile was stirred or
plowed 8" deep. The mixing was continued daily for
ten days with water being added throughout to maintain
about 10 percent moisture. After ten days, the layer of
75 ore which had been leached was repulped to 50 percent
6
The above description of the recovery of uranium
from low grade ores by the heap leaching process is
illustrative only and not restrictive. Obviously, the process
is equally effective for the recovery of metal values from
high grade ores either in small or large ore deposits. The
same heap leaching procedure was used to recover copper
from low grade copper oxide minerals and ,vanadium from
vanadium bearing sandstone and clay minerals. Sulfuric
acid was used as the leaching agent for these ores. The
process is not restricted to specific ores and leaching
agents but is equally applicable to any ore and leaching
agent. For example, the process can be applied to the
leaching of mixed copper oxide and copper sulfide ores
with acid ferric sulphate and acid ferric chloride, respectively.
Other examples of the application of this process
are: the leaching of manganese oxide ores with sulfurous
acid, the leaching of gold and silver ores with cyanide, the
leaching of tungsten ores with sodium carbonate and the
leaching of uranium ores with sodium carbonate.
'When acids are used as leaching agents, strong mineral
acids are preferred, such as, sulfurous, sulfuric, hydrochloric
and nitric acids. It is preferred to keep the acid
as concentrated as possible to provide a more concentrated
leaching solution; however, dilute acids can be
used.
In application of the process, the ore being treated
must be kept damp, but water should not be used in
amounts to make it sloppy. The leaching agent and the
moisture are, of course, plowed into the area 'of ore
which is being leached. The process can be applied
through successive top layers of a large pile of ore or
the ore can be piled on flat ground to the depth required
for a single leaching operation and the process performed
in this manner. The process can be applied to surface
35 mining to mine the ore in situ without removal from the
mine by ripping the ore and leaching in place.
The following examples of the process as applied to
recovery of copper and vanadium from low grade ores
are, again, illustrative but not restrictive of the process.
33.39
10.76
174.44 40
172.D7
136.62 45
180.77 50
Lb. U30 S
Feed to leaching, 62.30 tons at 0.140%
U30 S --------------------------
Feed to repulper, 76..46 tons at 0.1125%
UaOs1 -------------------------
Out:
Uranium recovered by ion exchange columns
(based on resin assays) 2 _
Lost to settled tailings (solids, 16.89;
solution, 16.50) _
Accountable accidental losses _
5
loads. of typical ore indicated preferable amounts of acid
of.68 and 83.5 lbs./ton and results obtained on the loads
from •which these samples were'taken indicated that results
did not vary appreciaJbly by adding more acid than
that indicated as adequate from the agitation leach tests. 5
Conversely, acid quantities far below that required in
the agitation leach tests were found to produce poor extractions.
For example, for another representative load,
agitation leach tests using acid at the rate of 170.5 lbs./
ton provided 77.9 percent extraction of uranium as U30s 10
and use of acid at a rate of only 60 lbs./ton resulted in
an extraction of only 32 percent.
Adequate cultivation is critical to obtaining good results.
This was illustrated in the leaching of load 5. The
load WllS spread out to the proper dimensions, acid added, 15
the heap stirred with the cultivator, moisture added, and
the heap cultivated again. During the next 17 days, moisture
was added to the heap but it was not cultivated. A
sample was then taken and the heap cultivated. The sample
taken before cultivation was washed and upon assay 20
was found to contain 0,02 percent U30 S' On the twentieth
day, after three days of cultivation, another sample was
taken, washed and assayed. This sample was found to
obtain only 0.007 percent U30 S' The cultivation uniformly
distributed the acid throughout the ore and resulted in 25
the lower uranium content of the tailings.
During pilot plant runs conducted over a three-month
period on the low grade uranium ore, an overall recovery
of 80.3 percent of the Ups in the ore into a finished
concentrate was obtained in processing 63.6 tons of ore 30
with an average feed grade of 0.14 percent U30 S' Acid
added for leaching averaged 79.5 lbs./ton (100% basis)
for the ores processed. The following table presents a
summary of the results obtained.
PILOT PLANT RESULtS
Overall metallurgical balance:
In:
Pilot plant recovery:
Based on calculated heads:
136.62
180.77_1O.76X100=80.3%
Based on assay heads:
136.62
174.44-10.76XI00=83.5%
Based on calculated heads and U30 S in elution solution:
121.26
165.41_1O.76 XIOO=78.5%
Reagent requirements:
Leaching: Lb./lb. U30 S 65
H~04 (100%) 39.5 lb./ton 16.70
Ion eHxchange (split elution technique):
2
S0
4
0.97
NaCI 7.71
U30S precipitation:
NH
a
---___________________________ 0.66
1 Additional weight in feed to repulper over that in feed
to leaching is due to picking up some of the earth underlying
the leaching pile.
2 U308 accounted for by assays of loaded elution liquor was
121.26 l!l.
3,777,004
8
mate amount by weight of leaching agent to ore required
for complete leaching of the ore: . . . .
9. The process of claim 1 III WhICh penodIc sample
tests are made of the tailings as the process progresses
to determine when leaching is complete.
5 10. The process of claim 1 in which a pile of the ore
after removal from the mine is made on the ground before
leaching having a height or depth equal to at least one
selected layer.
11. The process of claim 1 in which said selected layer
is from about 6 to about 12 inches in depth.
12. The process of claim 1 in which the ore is gold ore
and the leaching agent is a cyanide.
13. The process of claim 1 in which the ore is a silver
ore and the leaching agent is a cyanide.
14. The process of claim 1 in which the ore is tungsten
ore and the leaching agent is sodium carbonate.
15. The process of claim 1 in which the ore is uranium
ore and the leaching agent is sodium carbonate.
16. In the process for recovering metal values from
20 their ores by leaching the ore and recovering the metal
values from the resulting leach solution, the improvement
by which large lots of ore are leached.in situ ne~r the mine
site or in the mine without conventIonal eqUIpment and
25 housing which comprises breaking up the top layer of the
ore, leaching the top layer of the ore in situ by alternately
applying to the top layer until it is completely leached
water as necessary to keep it damp and leaching agent
accompanied by intermittent stirring of said top layer
30 separating said top layer of ore after it has been leached
for further processing to recover the metal value, and repeating
the process from the top of the lot downward on
successive layers until the complete lot of ore has been
leached.
35 17. A process for the recovery in situ of metal values
from a substantial lot of Ore which comprises:
(a) analyzing samples of the ore lot to determine the
approximate percentage by weight of leaching agent
to ore necessary for complete leaching;
(b) adding water uniformly to the top of a selected
depth of the lot of ore as necessary to insure that said
top layer is damp; .
(c) stirring said top layer of a selected depth to dIStribute
moisture uniformly therethrough;
(d) adding uniformly to the top of the layer a sufficient
amount of leaching agent as hidicated by said analysis
to leach the ore value therein;
(e) stirrmg said layer to distribute said leaching agent
uniformly therethrough;
(f) mtermittently stirring said layer and lldding water
as required to maintain the ore therein in a damp
condition over a period ranging from a few days to
several weeks until the metal value in said layer has
been substantially all leached as indicated by sample
analysis;
(g) removing said layer including the leaching solution
therein from the ore lot and forming a water slurry
of it;
(h) transferring the slurry to a settling basin and permitting
solids to settle out of the slurry;
(i) removing the clear solution from the settled solids
and recovering ore value from it; and
(j) repeating the above steps on successive layers of
the ore lot from the top down until all of the lot has
been leached.
18. The process of claim 17 in which the percentage of
water to ore is maintained at about 5.,..10%.
19. The process of claim 18 in which the leaching agent
is concentrated sulfuric acid.
70 20. The process of claim 19 in which the ore is uranium
ore and the metal value is uranium metal value.
21. The process of claim 19 in which the ore is.vanadiurn
ore and the metal value is vanadium metal value.
22. The process of claim 1 in which the ore is uranium
75 ore.
TABLE 3
7
solids with water filtered and washed. Vanadium values
were recovered from the leach solution by standard
procedures. The results obtained are tabulated as follows:
H
2
S0
4
, Ib.lton 400
Tails, percent V
2
0
S
1.53
Wt. loss, percent 18.2
Percent V20 S extraction ----------------------- 66.4 10
Lb. H2S04 per lb. V20 S extracted 8.9
The process has a distinct advantage when sUlfu~ic
acid is used to leach ore containing large limestone nbs
which are of no mineral value. Some ores have inclusions
of barren limestone incorporated therein. These occa- 15
sionally occur in sandstone uranium ores. When th~se ores
are leached with sulfuric acid in accordance wIth the
process of this invention a calcium sulfate coating is
formed on the large limestone particles which protects
them from acid, thus greatly decreasing the acid consumption
as compared to an agitatio.n le~ch pro~edure
wherein the limestone is in small partIcle SIze readlly attacked
by the acid. The process eliminates the use of oxidizing
agents, such as manganese dioxide, ordinarily required
in agitation leaching of some type ores. .
The process results in the generatlOn and .retent~on.of
heat in the ore being treated through the sl1ght dllutlon
of concentrated sulfuric acid when it is employed as a
leaching agent. Further, the process permits the use of
solar heat to assist the solubilizing action of the leaching
agent. Experience has shown that at least 80 percent of
the metal 'Values in low grade ores can be recovered by
the above process with a small amount of capital investment
with the overall expense being low enough to make
commercial operations feasible.
What is claimed is:
1. A process for the recovery in situ of metal values
from a substantial lot of ore which comprises:
(a) adding water and leaching agent for the metal value 40
to be recovered uniformly to the top layer of a selected
depth of the lot of ore in a sufficient amount to
leach the metal value in the selected depth;
(b) stirring said top layer of a selected depth to distribute
moisture and leaching agent uniformly therethrough;
45
(c) intermittently stirring said layer and adding water
as required to maintain the ore therein in a damp
condition over a period ranging from a few days to
several weeks;
(d) removing said layer including the leaching solution 50
therein from the ore lot and forming a water slurry
of it;
(e) transferring the slurry to a settling basin and permitting
solids to settle out of the slurry; .
(f) removing the clear solution from the settled sohds 55
and recovering metal value from it; and
(g) repeating the above steps on successive layers of
the ore lot from the top down until all of the lot
has been leached. 60
2. The process of claim 1 in which the leaching agent
is a strong mineral acid.
3. The process of claim 2 in which the leaching agent is
sulfuric acid.
4. The process of claim 3 in which the ore is a uranium 65
ore.
5. The process of claim 3 in which the ore is a vanadium
ore.
6. The process of claim 2 in which the acid is concentrated.
7. The process of claim 1 in which the moistu:e content
of the ore is maintained at about 10% of the weIght of the
ore.
8. The process of claim 1 in which s~mple tests m:e
made of the ore prior to step (a) to determme the apprOXI3,777,004
9
23. The process of claim 1 in which the ore is vanadium
ore.
24. The process of claim 1 in which the ore is gold ore.
25.. The process of claim 1 in which the ore is silver
ore. 5
26. The process of claim 1 in which the ore is tungsten
ore.
CARL D. QUARFORTH, Primary Examiner
15 R. L. TATE, Assistant Examiner
U.S. Cl. X.R.
75-101 R; 423-3,27,29,31,53,68
625,564
866,625
1,483,567
3,183,058
2,964,380
2,896,930
3,441,316
366,103
References Cited
UNITED STATES PATENTS
5/1899 Kendall 423-29
9/1907 Conedera 423-27
2/1924 i\njow 423-53
5/1965 Peter 23-321 X
1211960 Kolodney et al. 23=320
7/1959 Menke 23-320 X
4/1969 Hannifan et al. 75-101 X
7/1887 Hofmann 75-101
10
10
OTHER REFERENCES
De Andrade et al.: "Chemical Treatment of Uranium
Ores at the Mines in a Semi-Mobile Plant," 3rd Conf. on
Peaceful Uses, vol. 12, 1965, pp. 187-193.
Galkin et al.: Technology of Uranium, 1966, p. 103,
(AEC-tr-6638).
Application of Heap-Leaching to the Processing of Argentine
Ores, Cecchetto et aI., 3rd Conf. on Peaceful
Uses, vol. 12, 1965, p. 212.
Arden: Extraction and Refining of the Rarer Metals,
1957, pp. 130-1.
botto�(}01��0�eight: normal;mso-pagination:none;mso-layout-grid-align:none;text-autospace:none'>After the off-gas has been scrubbed of its loading of
30 dust and fume, it consists mainly of sulfur dioxide and
oxygen. The gas may be dried and compressed to liquefy
the sulfur dioxide. The oxygen remains in the gaseous
form and is recycled to the flash roaster.
Table 3 shows the results from re-roasting a number
of calcines:
After re-roasting of the calcine, portions of the copper,
the remaining rhenium not collected in the scrubbers,
and a portion of the sulfur are soluble in mineral
50 acid solutions. Since the process produces a dilute mineral
acid -sulfurous acid- in the scrubbers, it is used
to leach the copper and the remaining rhenium from
the calcine (Table 4).
55 TABLE 4.-REMOYAL OF COPPER AND RESIDUAL RHENIUM
AND SULFUR FROM RE-ROASTED CALCINES
Molybdenum balance BY LEACHING WITH SULFUROUS ACID
(percent distribution)
88 94 6
84 94 6
76 ..
95 .
125
140
218
40
Rhenium balance
567
707
736
736
TABLE I.-ROASTING TEST RESULTS
Preheat temperatures: 650-750'C. range
Hearth temperature: 550-650'C. range
Percent stoichiometric oxygen: 170-240
the latter being the only other gaseous component in
the exhaust stream· pertinent to this control feature.
The sulfur dioxide-oxygen ratio in the exhaust stream
can be partially controlled to provide the optimum
value, if necessary, by the introduction of sulfur dioxide 5
gas with the oxygen. The ratios reflected by 30-35%
volume of sulfur dioxide are by no means critical, but
its use to provide favorable reaction zone conditions
illustrates the effectiveness of this method of control.
There are three other principal parameters affecting
the temperature control and/or the oxidationvolatilization
process, one or all of which may be used
to control these factors in varying degrees. These parameters
are: (1) preheat temperature, (2) the height
ofthe reactor column, and (3) heat dissipation from 15
the column. The first of these, like the oxygen-sulfur
dioxide ratio, is applied during the operation of the process..
The latter two are built-in to the construction of
the apparatus.
The preheat temperature is readily controlled by ad- 20
justing the heat input to the indirect-fired preheat furnace.
The height of the reactor column determines the
dwell time of the sulfide particles in the reaction zone
for complete oxidation and for formation and volatil- 25
ization of rhenium oxide. The optimum height for a
given operation is developed by calculations and measurements
derived from pilot plant operation. For example,
in a continuous pilot plant operation excellent
results were obtained using a vertical column 44 inches
in height and 6 inches in diameter with a rotating
hearth 3 feet in diameter. These dimensional relationships
are not critical and would change with change in
other variables, such as, concentrate characteristics,
composition of feed gases, rate of gas injection, etc. 35
The heat dissipated from the vertical column is controlled
by design, and construction materials used. The
construction can be varied from highly insulated construction
to high conductivity construction with a cooling
media. The radiation and convection loss of heat 40
generated for a metal conducting material and a given
feed rate can be readily calculated. Additional heat
may be removed from the column by water cooling or
other heat exchange media.
The results given below are illustrative of those obc 45
tained by application of the above-described process in
conjunction with the apparatus described.
Table 1 shows some material balances obtained in
roasting tests performed on molybdenite concentrate.
1.. ..
2 ..
3 ..
4 .
Rhenium Product Yolatil- Dust and
T_es_t ___--f,e-ed_--,(p-p_m)--,iz-ed_(_%)__P_ro_du_ct__sc_ru_bb_er 60 Spiaem-
----------------------
The data in Table 1 shows the variability of the rhenium
content of the product produced at somewhat
65 About 7 percent of the molybdenum contained in
the calcines is also solubilized in the sulfurous acid
leach.
3,770,414
JlO
zone,
c. controlling the temperature in the first oxidation
zone durin~ the introduction thereto of said preheated
particles and thereafter to maintain a temperature
therein above the volatilization temperature
of rhenium oxide and below the volatilization
temperature of molybdic oxide to form rhenium
oxide, sulfur dioxide and molybdic oxide, which
latter oxide along with other solids passes to a second
oxidation zone where any unoxidized molybdenite
is completely oxidized, said second oxidation
zone being hec;ted by exothermic heat of the
reactions occurring therein,
d. passing oxygen through said second oxidation zone
to oxidize molybdenite contained therein,
e. passing at least some of the oxygen travelling to
said first oxidation zone through said second oxida,
tion zone to heat the oxygen before it reaches the
first oxidation zone,
20 f. recovering rhenium oxide by collecting it in a recovery
zone outside the first oxidation zone and
dissolving it in water,
g. recovering rhenium from the water solution of rhenium
oxide, and
25 h. recovering insoluble molybdic oxide from the second
oxidation zone.
2. The process of claim 1 in whi'ch said concentrate
is preheated to about 500°C.
30 3. The process of claim 1 in which molybdenum values
are recovered.
4. The process in claim 1 in which rhenium values are
recovered.
5. The method of claim 1I in which the temperature
of the first oxidation zone resulting from exothermic
heat of reaction is controlled by controlling the reaction
rate of the oxidation reactions occurring therein.
6. The method of claim 5 in which said reaction rate
is controlled by adjusting the relative feed rate of oxy40
gen and molybdenite concentrate to the first oxidation
zone to control the stoichiometric ratio of oxygen to
metal sulfides introduced therein.
7. The method of claim 6 in which said stoichiometric
ratio is at least one.
S. The method of claim If) in which said stoichiometric
ratio is at least 120%.
9. The method of claim 6 in which sulfur dioxide is
introduced to the first oxidation zone.
10. The method of claim 6 in which the sulfur diox50
ide-oxygen ratio in the exhaust gases from the first oxidation
zone is used to determine the relative rate of addition
of oxygen and concentrate.
1I1. The method of claim 1 in which the exhaust gas
contains up to about 50% by volume of sulfur diOldde.
55 12. The method of claim 1 in which the dwell time of
concentrate particles in the first oxidation zone is controlled
by varying the diameter and height of said zone.
1I3. The process of claim 1I in which oxygen in the exhaust
gases is recycled for reuse in the method.
60 14. The process of claim 1 in which sulfur dioxide in
the exhaust gases is dissolved in water to form sulfurous
acid and the sulfurous acid used to leach impurities
from the molybdic oxide calcine recovered from the
second oxidation zone.
* >I< >I< * *
About 7% of the molybdenum contained in the calcines
is also solubilized in the sulfurous acid leach.
The leached residue is separated from the leach solution
by filtration and after drying is ready for packaging 5
for sale. The leach solution joins the solutions from the
scrubbers on the flash roaster and re-roaster.
The effectiveness of the above-described process is
graphically illustrated by the high recovery of rhenium
and molybdenum achieved. it provides for the recovery 10
of up to 95% of rhenium and high recovery ofmolybdenum
in molybdenite with a minimum of process time
and a minimum of oxygen and added heat. The economic
advantages of these features are apparent. The
process is adaptable to either a batch or continuous op- is
eration.
It is an attractive side advantage of the. process that
a small volume of exhaust gas containing a high percentage
by volume of sulfur dioxide is produced. The
process is normally operated with an exhaust gas volume
discharge rate of 1,350 cubic feet per minute
(CFM) with up to 220% excess oxygen and 30-50% by
volume of sulfur dioxide in the exhaust gas. This high
volume percentage of sulfur dioxide makes its recovery
economically feasible for various commercial uses. In
contrast, present-day processes utilizing air for cooling
and for supplying oxygen are of necessity operated with
an exhaust volume discharge rate of 40,000 CFM, 16
volume percent excess oxygen and 1-2 volume percent
of sulfur dioxide. This volume percentage of sulfur dioxide
in the exhaust gas is so low that its recovery is not
economically feasible because it involves processing
such large volumes of gas. As a result the sulfur dioxide
is exhausted to the atmosphere creating a serious pollution
problem in heavily populated areas. The process of 35
this invention eliminates this problem.
The reduced volume of exhaust gas also results in a
much higher concentration of rhenium oxide in the exhaust
gas than is obtained in conventional processes.
As a result, recovery of substantially all of the rhenium
is far more feasible and economical than in present processes
using air with resultant large volumes of exhaust
gas to be processed for recovery of the rhenium oxide.
Reduction of the volume of gas processed through
the system by a factor of about 30resultsin a dr1!§jic: 45
reduction in the size of equipment require-d~ith~jgnificant
savings in equipment cost and floor space.
What is claimed is:
n. A method for recovering rhenium and molybdic
mdde from molybdenite concentrate which comprises:
a. pre-heating particles of said concentrate in an oxygen-
free atmosphere to a temperature not in excess
of about 750"C to raise the temperature of the particles
to promote flash oxidation of the molybdenite
when the particles are introduced into a flash
oxidation zone,
b. causing said pre-heated particles to fall through a
first oxidizing zone of heated oxygen with said particles
and heated oxygen moving countercurrent to
each other to disperse said pre-heated molybdenite
particles in said heated oxygen to provide maximum
particle surface contact with heated oxygen
for effective oxidation, said first oxidation zone
being heated substantially by the exothermic heat
of the reactions occurring in said first oxidation 65